EP3149214B1 - Traitement hydrométallurgique de boues anodiques - Google Patents

Traitement hydrométallurgique de boues anodiques Download PDF

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Publication number
EP3149214B1
EP3149214B1 EP15730210.0A EP15730210A EP3149214B1 EP 3149214 B1 EP3149214 B1 EP 3149214B1 EP 15730210 A EP15730210 A EP 15730210A EP 3149214 B1 EP3149214 B1 EP 3149214B1
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Prior art keywords
leaching
filtrate
accomplished
residue
selenium
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EP15730210.0A
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German (de)
English (en)
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EP3149214A1 (fr
Inventor
Henri Virtanen
Sönke Schmachtel
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Outotec Finland Oy
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Outotec Finland Oy
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Priority to RS20191430A priority Critical patent/RS59570B1/sr
Priority to PL15730210T priority patent/PL3149214T3/pl
Publication of EP3149214A1 publication Critical patent/EP3149214A1/fr
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    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B11/00—Obtaining noble metals
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B11/00—Obtaining noble metals
    • C22B11/04—Obtaining noble metals by wet processes
    • C22B11/042—Recovery of noble metals from waste materials
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
    • C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
    • C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
    • C22B3/08—Sulfuric acid, other sulfurated acids or salts thereof
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
    • C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
    • C22B3/10—Hydrochloric acid, other halogenated acids or salts thereof
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
    • C22B3/44—Treatment or purification of solutions, e.g. obtained by leaching by chemical processes
    • C22B3/46—Treatment or purification of solutions, e.g. obtained by leaching by chemical processes by substitution, e.g. by cementation
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B61/00—Obtaining metals not elsewhere provided for in this subclass
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
    • C22B7/006—Wet processes
    • C22B7/007—Wet processes by acid leaching
    • Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
    • Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
    • Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
    • Y02P10/00—Technologies related to metal processing
    • Y02P10/20—Recycling

Definitions

  • the present invention relates to hydrometallurgical treatment of anode sludge, and more particularly to a method of separating precious metals from anode sludge obtained from copper electrolysis.
  • hydrometallurgical methods for separating the precious metals from anode sludge are based on the use of nitric acid, because the solubility of silver as nitrate is high.
  • hydrometallurgical processes based on the use of nitrates for treating anode sludge are not compatible with the rest of the electrolytic process, because the electrolytic refining of copper is carried out in a sulfate solution.
  • the nitrate bearing sludge must be mechanically ground finer, in order to make the leaching successful.
  • WO2005028686 is known a method for processing anode sludge obtained from copper electrolysis. Said method includes at least an atmospheric leaching step as well as a step of leaching into a neutral aqueous solution.
  • the leaching residue obtained from the aqueous leaching can optionally be leached further into sulfuric acid, and the thus obtained residue can subsequently be leached into hydrochloric acid.
  • Selenium is distilled from the obtained, precious metals bearing precipitate, and the distilling residue is returned to the process or processed outside the plant.
  • the leaching residue from wet chlorination is processed further for recovering the lead and silver contained therein. After lead separation, the silver chloride from the precipitate is leached into an ammonia solution, reprecipitated as pure chloride and finally reduced into metallic silver.
  • An object of the present invention is thus to provide a method so as to overcome the above problems.
  • the objects of the invention are achieved by a method and an arrangement, which are characterized by what is stated in the independent claims.
  • the preferred embodiments of the invention are disclosed in the dependent claims.
  • the invention is based on the idea of a straightforward hydrometallurgical processing of anode sludge where no selenium calcination or slimes smelting is needed.
  • An advantageous feature of the method and arrangement of the present invention is that the new method makes use of such chemicals, for example sulfuric acid, that are normally used in an electrolytic copper refinery. This enables the recirculation of the solutions to the electrolysis or to the electrolyte purification. A remarkable reduction is achieved as regards to environmental emissions, because harmful gas emissions released from selenium calcination and slimes smelting are avoided. The total process delay is also cut.
  • chemicals for example sulfuric acid
  • FIG. 1 shows a flowchart of a first embodiment of the method of the present invention.
  • the present invention provides a method of separating precious metals from anode sludge obtained from copper electrolysis.
  • the method of separating precious metals from anode sludge obtained from copper electrolysis comprises
  • step (c) Typically the method further comprises, prior to step (c),
  • FIG 1 shows a flowchart a first embodiment of separating precious metals from anode sludge obtained from copper electrolysis.
  • anode sludge 1 obtained from the electrolytic refining of copper is introduced into a first leaching step (a) 101, wherein leachable chlorides are removed from the anode sludge.
  • the anode sludge is a mixture of different compounds comprising copper, selenium, precious metals and as impurities other metals and elements, such as lead.
  • the anode sludge is obtained from the electrolytic refining of copper, and the composition of said raw material can vary.
  • the copper content of the raw sludge can be over 30%.
  • the silver and selenium content of such sludge is typically about 5 to 10%, and its impurity contents (As, Sb, Bi, Pb, Te, Ni) are typically of the order of 1 to 5%.
  • Leaching in step (a) is accomplished in an aqueous sulfuric acid solution 10.
  • the solution preferably contains from 50 to 150 g/L H 2 SO 4 .
  • the leaching step (a) is accomplished under an elevated temperature, typically from 60 to 95°C, preferably from 80 to 90°C.
  • the slurry density is typically less than 500 g/L, preferably from 200 to 300 g/L.
  • Leaching in step (a) is advantageously performed under atmospheric pressure.
  • Advantageously leaching in step (a) is performed under oxidizing conditions (11) for oxidizing copper present in the anode sludge, but can also be performed without any oxidizing.
  • Leaching of the anode sludge in an aqueous sulfuric acid solution in step (a) is preformed to remove leachable chlorides from the anode sludge.
  • the removal of leachable chlorides in step (a) reduces precipitation of silver chloride in further process steps and thus increases yield in silver recovery.
  • some of the copper and arsenic comprised in the anode sludge are leached in this step.
  • After filtration a chloride containing filtrate 19 and a first leaching residue 2 depleted of leachable chlorides is obtained.
  • the first leaching residue 2 is then introduced into a pressure leaching step (b) 102.
  • the first leaching residue 2 obtained from step (a) and entering into the pressure leaching step (b) 102 is typically diluted to a slurry density of less than 300 g/L, preferably from 100 to 200 g/L depending on the silver concentration of the sludge.
  • Leaching in step (b) is accomplished in an aqueous sulfuric acid solution 12.
  • the leaching solution preferably contains from 200 to 500 g/L H 2 SO 4 .
  • the pressure leaching step (b) is accomplished under an elevated temperature, typically more than 140°C, preferably from 160 to 170°C. Leaching is advantageously performed under an elevated pressure from 5 to 9 bar. In this step silver and selenium are solubilized and some impurities such as tellurium, arsenic, copper and nickel can be leached away.
  • An oxidizing agent 13 is used in the pressure leaching step (b) 102 to improve dissolution of silver and selenium.
  • the oxidizing agent 13 is oxygen or hydrogen peroxide, preferably oxygen.
  • the pressure leaching step (b) can be repeated one or more times, preferably one time, to ensure dissolution of silver and selenium.
  • a first filtrate 20 comprising silver and selenium and a second leaching residue 3 mostly depleted of silver and selenium is obtained.
  • the first filtrate 20 typically also comprises tellurium.
  • the obtained second leaching residue 3 is then introduced into either directly to leaching step (c) 105 or preferably to an optional leaching step (d) 103 and/or optional leaching step (e) 104.
  • the method of the present invention further comprises recovering silver and, optionally, selenium from the first filtrate 20.
  • the filtered solution obtained from step (a) 20, and optionally also the filtered solution obtained from bismuth recovery 31, as discussed below, are introduced into a silver and selenium recovery step 112.
  • the recovery of silver and selenium is accomplished by chloride precipitation and sulfur dioxide cementation.
  • Silver is first precipitated by adding stoichiometric amount of hydrochloric acid.
  • Metallic silver can be produced from silver chloride by known methods e.g. silver oxide precipitation followed by silver reduction. After silver precipitation selenium can be precipitated as elementary selenium by reduction with sulfur dioxide (70).
  • the filtrate 21 from the recovery of silver and selenium 112 can be further treated to recover tellurium from the filtrate in a tellurium recovery 122.
  • Tellurium is preferably recovered from the filtrate by cementing with copper 71 into Cu 2 Te 61.
  • the remaining solution 22 can be further treated in the electrolyte purification of copper electrolysis.
  • the recovery of silver and selenium can be accomplished for example by copper cementation with e.g. copper powder or copper chips or precipitation as silver selenide by using sulfur dioxide.
  • the second leaching residue 3 can be subjected to an optional leaching step (d) and/or to an optional leaching step (e).
  • Leaching in step (d) is accomplished in an aqueous sulfuric acid solution 14.
  • the leaching solution preferably contains from 400 to 900 g/L H 2 SO 4 .
  • the slurry density in step (d) is less than 400 g/L, preferably from 200 to 300 g/L.
  • the leaching step (d) is accomplished under an elevated temperature, typically from 80 to 120°C, preferably from 90 to 110°C.
  • Leaching in step (d) is advantageously performed under atmospheric pressure.
  • Leaching step (d) mainly provides for dissolution of bismuth (113) from the second leaching residue. However, also any remaining silver, selenium and/or tellurium will be leached.
  • a third filtrate 30 and a further treated second leaching residue 4 depleted of Bi is obtained.
  • the further treated second leaching residue 4 is then introduced either directly into a leaching step (c) 105 or preferably into a second optional leaching step (e) 104.
  • the optional leaching step (e) 104 provides for removal of remaining impurities before silver and PGMs separation in step (c) 105.
  • Leaching in step (e) is performed in an aqueous hydrochloric acid solution to obtain a fourth filtrate 40 comprising Pb, and optionally Se and Te, and a still further treated second reaching residue 5 further depleted of lead and any remaining Se and Te.
  • Leaching in step (e) is performed in the absence of an oxidant to ensure that gold and PGMs are not solubilized.
  • the non-oxidative leaching in an aqueous solution of hydrochloric acid 15 is mainly utilized to dissolve lead. However, significant amount of any remaining selenium is also dissolved. Also remaining tellurium present in the second leaching residue 3 or in the further depleted second leaching residue 4 will be dissolved. Rest of the selenium is dissolved in oxidative hydrochloric acid leaching step (c) where main aim is to leach gold and PGMs.
  • a fourth filtrate 40 and a still further treated second leaching residue 5 depleted of lead is obtained.
  • the still further treated second leaching residue 5 is then introduced to a leaching step (c) 105.
  • the method of the present invention additionally comprises recovering of lead from the fourth filtrate 40. Accordingly the filtered solution 40 is obtained from step (e) is introduced into a lead recovery step 114. Lead chloride 63 can be crystallized by temperature decrease.
  • the filtered solution 41 is introduced into a selenium and/or tellurium recovery step 124.
  • Selenium and/or tellurium can be recovered e.g. by reduction with sulfur dioxide 72 as discussed above.
  • the filtered solution 42 obtained from the recovery of selenium and/or tellurium 64 in step 124 can be further subjected to a neutralization step 125.
  • the second treated leaching residue 3, the further treated second leaching residue 4, or the still further treated second leaching residue 5 is leached in an aqueous hydrochloric acid solution 16 to dissolve gold and platinum-group metals (PGMs) and to obtain a second filtrate 50 comprising Au and PGMs and a final leaching residue 6.
  • the solution preferably contains from 150 to 250 g/L HCI.
  • the slurry density in step (c) is less than 500 g/L, preferably from 200 to 400 g/L.
  • Leaching step (c) is typically accomplished under an elevated temperature, typically form 70 to 90°C, preferably from 75 to 85°C. Leaching in step (c) is advantageously performed under atmospheric pressure.
  • An oxidizing agent 17 is used in the leaching step (c) to improve dissolution of gold and PGMs.
  • the oxidizing agent 17 is hydrogen peroxide or chlorine, preferably hydrogen peroxide.
  • the method of the present invention further comprises recovering of Au and, optionally, PGMs from the second filtrate 50.
  • the filtered solution 50 obtained from step (c) is introduced into the gold and PGMs recovery step 115.
  • the recovery of gold is accomplished by reduction of gold e.g. by sulfur dioxide or ferrous chloride.
  • Gold is advantageously reduced by treating the filtrate 50 with SO2 gas, so that gold is precipitated (65) in two steps. In the first step, pure gold is precipitated.
  • the impure gold obtained from the second step can be recycled back to the leaching step (c).
  • gold can be recovered by solvent extraction e.g. using dibutyl carbitol as extractant. From the extraction solution gold can be directly reduced to gold powder. Precipitation of gold by SO2 is preferred as it is more economic and simpler method for recovering gold.
  • PGMs can be recovered (65).
  • the recovery of platinum-group metals is accomplished by cementation of PGMs.
  • PGMs can be cemented by iron to obtain a mixture containing platinum-group metals.
  • the filtrate 51 obtained from the recovery of Ag and PGMs is then typically subjected to a neutralization step 125, optionally together with the filtrate 42 obtained from the recovery of selenium and/or tellurium in step 124.
  • anode sludge was treated by the method of the present invention to leach and recover silver and selenium.
  • Sludge containing 7.9% Ag, 7.9% Se, 1.9% Te, 23.1% Cu, 4.8 As, and 0.47% CI was first leached in a sulfuric acid solution containing 100 g H 2 SO 4 /l at a temperature of 90°C in a four liters reactor. Solids concentration in the start of leaching was 250 g/l and duration of leaching was 2 hours. After leaching residue was filtered, washed and analyzed together with the filtrate. Residue cake concentrations were 9.1% Ag, 8.8% Se, 1.8% Te, 13.2% Cu, 2.5 As, and 0.07% CI.
  • Filtrate contained 36.5 g Cu/L, 6.8 g As/L, 1.2 g Cl/L, 0.6 g Te/L, and 0.1 g Se/L and recovery of chloride was about 80%. Practically no silver was found from the solution.
  • Silver from the filtrate was precipitated as silver chloride by stoichiometric hydrochloric acid addition. After silver chloride precipitation selenium was precipitated at elevated temperature by sulfur dioxide and finally tellurium was cemented by using copper powder.
  • Sludge after pressure leaching contained 5.3% bismuth.
  • For bismuth dissolution sludge was leached in sulfuric acid solution having 800 g H 2 SO 4 /L and solids concentration 250 g/L. Temperature during leaching was kept slightly above 100°C. After two hours leaching period slurry was filtered hot to prevent problems due to crystallization of bismuth sulfate (62) during filtration. Final bismuth concentration in the filtrate at the leaching temperature was 6.4 g/L. Also some silver, selenium, tellurium and arsenic were leached at this stage and washed residue contained 0.8% Ag, 0.4% Se, 0.6% Te, 2.2% As, and 2.1% Bi. Achieved bismuth concentration in the filtrate tells that by lowering solids concentration in the start even more bismuth can be leached.
  • Recoveries over the three leaching stages were 96.2% for silver, 98.1% for selenium, 89.1% for tellurium, 82.9% for arsenic, and 70.9% for bismuth based on analyses, weights and volumes of residues and filtrates.
  • Residue from pressure leaching containing 1.1% Ag, 0.3% Se, 0.6% Te, 0.5% As, 2.9% Sb, 1.9% Bi, and 4.8% Pb was first leached in hydrochloric acid solution containing 100 g HCI/L at a temperature of 85°C. Aim of the experiment was to leach lead and solids concentration was 100 g/L because of the limited lead chloride solubility. Duration of leaching was three hours and no oxidant was used. After leaching slurry was filtered, cake washed and analyzed together with filtrate. Residue cake contained 1.2% Ag, 0.14% Se, 0.5% Te, 0.5% As, 2.5% Sb, 1.4% Bi, and 0.4% Pb.
  • Lead concentration in the filtrate was 4.2 g/L and selenium and tellurium concentrations 0.12 and 0.14 g/L. Analyzed gold concentration was 30 mg/L. Selenium, tellurium and gold were precipitated from filtrate by sulfur dioxide before cooling of the solution and lead chloride crystallization.

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Claims (15)

  1. Procédé séparation de métaux précieux d'une boue anodique obtenue à partir de l'électrolyse du cuivre, comprenant :
    (a) la lixiviation de la boue anodique dans une solution aqueuse d'acide sulfurique pour éliminer les chlorures lixiviables et obtenir un premier résidu de lixiviation appauvri en chlorures ;
    (b) la lixiviation sous pression du premier résidu de lixiviation pour dissoudre Ag et Se et obtenir un premier filtrat comprenant Ag et Se et un second résidu de lixiviation appauvri en Ag et Se ;
    (c) la lixiviation du second résidu de lixiviation avec une solution aqueuse d'acide chlorhydrique pour dissoudre Au et les métaux du groupe platine (PGM) afin d'obtenir un deuxième filtrat comprenant Au et les PGM et un résidu de lixiviation final.
  2. Procédé selon la revendication 1, dans lequel, avant l'étape (c), le procédé comprend :
    (d) la lixiviation du second résidu de lixiviation obtenu à l'étape (b) dans une solution aqueuse d'acide sulfurique pour dissoudre le bismuth et l'argent, le sélénium et/ou le tellure restant éventuellement pour obtenir un troisième filtrat comprenant Bi et éventuellement Ag, Se et/ou Te et un autre second résidu de lixiviation traité appauvri en outre en au moins Bi ; et/ou
    (e) la lixiviation du second résidu de lixiviation obtenu à l'étape (b) ou (d) dans une solution aqueuse d'acide chlorhydrique pour dissoudre le plomb et le sélénium et/ou le tellure restant éventuellement afin d'obtenir un quatrième filtrat comprenant Pb et éventuellement Se et/ou Te et un autre second résidu de lixiviation traité encore appauvri en outre en au moins le plomb.
  3. Procédé selon la revendication 1 ou 2, dans lequel la lixiviation à l'étape (a) est effectuée dans des conditions oxydantes pour oxyder le cuivre présent dans la boue anodique.
  4. Procédé selon l'une quelconque des revendications 1 à 3, comprenant en outre la récupération d'Ag et Se du premier filtrat.
  5. Procédé selon la revendication 4, dans lequel la récupération d'Ag et Se est effectuée par cémentation au dioxyde de soufre.
  6. Procédé selon la revendication 4, dans lequel la récupération d'Ag est effectuée par précipitation au chlorure d'argent.
  7. Procédé selon l'une quelconque des revendications 1 à 6, comprenant en outre la récupération d'Au et des PGM du deuxième filtrat.
  8. Procédé selon la revendication 7, dans lequel la récupération d'Au est effectuée par réduction de l'or.
  9. Procédé selon la revendication 7 ou 8, dans lequel la récupération des PGM est effectuée par cémentation des PGM.
  10. Procédé selon l'une quelconque des revendications 1 à 9, dans lequel l'étape de lixiviation (a) est effectuée à température élevée, typiquement de 60 à 95 °C, de préférence de 80 à 90 °C, et à pression atmosphérique.
  11. Procédé selon l'une quelconque des revendications 1 à 10, dans lequel l'étape de lixiviation (b) est effectuée à température élevée, typiquement de plus de 140 °C, de préférence de 160 à 170 °C, et à pression élevée de 5 à 9 bars.
  12. Procédé selon l'une quelconque des revendications 1 à 11, dans lequel l'étape de lixiviation (c) est effectuée à température élevée, typiquement de 70 à 90 °C, de préférence de 75 à 85 °C, et à pression atmosphérique.
  13. Procédé selon l'une quelconque des revendications 1 à 12, dans lequel un agent oxydant, de préférence de l'oxygène ou du peroxyde d'hydrogène, mieux encore de l'oxygène, est utilisé dans l'étape de lixiviation sous pression (b).
  14. Procédé selon l'une quelconque des revendications 1 à 13, dans lequel la lixiviation à l'étape (b) est effectuée dans une solution aqueuse d'acide sulfurique contenant de préférence 200 à 500 g/L de H2SO4.
  15. Procédé selon l'une quelconque des revendications 1 à 14, dans lequel un agent oxydant, de préférence du peroxyde d'hydrogène ou du chlore, mieux encore du peroxyde d'hydrogène, est utilisé dans l'étape de lixiviation (c).
EP15730210.0A 2014-05-28 2015-05-27 Traitement hydrométallurgique de boues anodiques Active EP3149214B1 (fr)

Priority Applications (2)

Application Number Priority Date Filing Date Title
RS20191430A RS59570B1 (sr) 2014-05-28 2015-05-27 Hidrometalurška obrada anodnog mulja
PL15730210T PL3149214T3 (pl) 2014-05-28 2015-05-27 Hydrometalurgiczna obróbka szlamu anodowego

Applications Claiming Priority (2)

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FI20145484A FI126064B (en) 2014-05-28 2014-05-28 Hydrometallurgical treatment of anode sludge
PCT/FI2015/050366 WO2015181446A1 (fr) 2014-05-28 2015-05-27 Traitement hydrométallurgique de boues anodiques

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EP3149214A1 EP3149214A1 (fr) 2017-04-05
EP3149214B1 true EP3149214B1 (fr) 2019-08-14

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US (1) US10316389B2 (fr)
EP (1) EP3149214B1 (fr)
AU (1) AU2015265793B2 (fr)
CA (1) CA2949916C (fr)
CL (1) CL2016003004A1 (fr)
ES (1) ES2751598T3 (fr)
FI (1) FI126064B (fr)
PL (1) PL3149214T3 (fr)
RS (1) RS59570B1 (fr)
RU (1) RU2650663C1 (fr)
WO (1) WO2015181446A1 (fr)

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US11408053B2 (en) 2015-04-21 2022-08-09 Excir Works Corp. Methods for selective leaching and extraction of precious metals in organic solvents
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EP3149214A1 (fr) 2017-04-05
RU2650663C1 (ru) 2018-04-16
FI126064B (en) 2016-06-15
CL2016003004A1 (es) 2017-05-05
FI20145484A (fi) 2015-12-23
AU2015265793B2 (en) 2018-02-22
ES2751598T3 (es) 2020-04-01
PL3149214T3 (pl) 2020-05-18
WO2015181446A1 (fr) 2015-12-03
US20170198370A1 (en) 2017-07-13
RS59570B1 (sr) 2019-12-31
CA2949916C (fr) 2022-07-12
AU2015265793A1 (en) 2016-12-22
US10316389B2 (en) 2019-06-11
CA2949916A1 (fr) 2015-12-03

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